7.1 Parameters of working face and equipment
At the No. 8105 LLTCC coalface in Tashan coal mine, the face length is 207 m, the maximum cutting height is 5.0 m, and the cutting depth is 0.8 m. The installed capacity for the working face is 10 Mtpa. Based on the technical parameters and performance of fully mechanized coal mining equipment in China, the main technical parameters of the fully mechanized coalface equipment are determined as shown in Table 27.
In April 2010, the combined trial operation on surface of the complete equipment developed was conducted at Tashan coalmine (Fig.71). From October to December in the same year, an underground industrial trial was completed.
7.2 Operation of trial equipment and production
During the period of the industrial trial at the No. 8105 face from October to December 2010, the statistics of produc- tion downtime caused by major equipment accidents are shown in Table 28. From October to December 2010, the
Table 27 Equipment for the coalface
No. Equipment name Main parameters of equipment Reference type
1 Hydraulic support Height:2800–5200 mm; center distance: 1750 mm; support intensity: [1.35 MPa ZF15000/28/52
2 Shearer Cutting power: 2 9 750 kW; cutting height: 2.7–5.0 m; cutting depth: 800 mm; AC electric
traction; voltage: 3300 V
MG750/1915-GWD
3 Front scraper
conveyor
Length: 220 m; power: 2 9 855 kW; voltage: 3300 V; transportation capacity: 2500 t/h SGZ1000/2 9 855
4 Rear scraper
conveyor
Length:220 m; power:2 9 1000 kW; voltage: 3300 V; transportation capacity: 3000 t/h SGZ1200/2 9 1000
4 Stage loader Transportation capacity: 3500 t/h; power: 450 kW; voltage: 3300 V PF6/1542
5 Crasher Crashing capacity: 4250 t/h; power: 400 kW;voltage: 3300 V SK1118
6 Emulsion pump Working pressure: 31.5 MPa, Rated flow rate: 400 L/min; motor power: 250 kW,
voltage: 3300 V, fluid box: 2, volume: 2500 L
BRW400/31.5
7 Spray pump Working pressure: 12.5 MPa, Rated flow rate: 500 L/min,motor power: 132 kW,
voltage: 3300 V, fluid box: 1,volume: 3000 L
BPW500/12.5
Table 28 Statistics of production downtime caused by major equipment failure during the industrial trial period from October to December
2010
Time Shearer Hydraulic support Scraper conveyor Belt conveyor Power supply system
October 2010 22 h 35 min 0 0 8 h 30 min 2 h 35 min
November 2010 4 h 0 15 h 50 min 5 h 10 min 4 h
December 2010 4 h 30 min 0 1 h 25 min 15 min 0
Total 31 h 5 min 0 17 h 15 min 13 h 55 min 6 h 35 min
Fig. 71 Combined trial operation on the surface of complete
production downtime caused by shearer accidents was 31 h and 5 min, the shearer usage rate was 98.45 %, and its performance was stable. The production downtime caused by hydraulic support failure was zero, its working resis- tance varied within a 9000–13,000 kN range, and overall the support was in good condition. The production down- time caused by scraper conveyor failure was 17 h and 15 min with an usage rate of 98.95 %. The production downtime caused by belt conveyor accidents was 13 h and 55 min, with an usage rate of 99.15 %. The production downtime caused by power system accident was 6 h and 35 min, failure-free rate was 99.61 %.
In October 2010, the production downtime caused by cumulative electrical accidents during the production per- iod was about 57 h and 40 min. Based on four shifts with six hours each shift in the working face, the usage rate was 89.3 %. In November 2010, the production downtime caused by cumulative electrical accidents during the pro- duction period was about 53 h and 50 min, the usage rate was 91.3 %. In December 2010, the production downtime caused by cumulative electrical accidents during the pro- duction period was about 7 h and 10 min, the usage rate was 95.7 %. In the industrial test period from October 2010 to December 2011, average usage rate of equipment in use was 92.1 %.
During the industrial test period from October to December 2010 at No. 8105 coalface, the cumulative face advance was 472.5 m, average daily advance was 5 m,
Table 29 Production statistics at No. 8105 coalface during the industrial test period from October to December 2010 Time Number of working days Cutting height average/ maximum (m) Face advance (m) Production (t)
Average/maximum monthly production
(t) Average daily production (t) Average manpower efficiency (t/worker) Coal recovery (%) Equipment usage (%) October 2010 30 4.5/5.0 147.75 901,732 907527/916785 30057.8 362.2 88.9 92.1 November 2010 30 152.25 904,065 30135.5 363 December 2010 30 172.5 916,785 30559.5 368.2 Total 90 4.5/5.0 472.5 2,722,582 907527/916785 30250.9 364.5 88.9 92.1
Fig. 72 Displacement curve of the roadway surface. a During
maximum daily advance was 8 m; the cumulative coal production was 2.723 Mt, the average daily production was 30,300 t, maximum daily production was 38,000 t; average manpower efficiency was 364.5 t/worker. In three con- secutive months during the industrial test, average monthly production reached more than 900,000 t, coal recovery rate at the working face reached 89.9 %. In 2011, the average monthly coal production was 904,000 t, the cumulative coal production for the year was 10.849 mt, achieving the goal of producing 10 Mtpa. The production statistics for the industrial test at the No. 8105 coalface from October to December 2010 is listed in Table29.
7.3 Analysis of roadway support
During the excavation and support in the No. 5105 road- way, the deformation of the surrounding rock of the roadway was monitored, see Fig.72a. The wall-to-wall convergence and the roof-to-floor convergence during the roadway excavation period was 25 mm and 16.6 mm, respectively. The support effect during the excavation period is shown in Fig.73. In general, the deformation was small during roadway excavation period, and the effect of roadway support was good.
After the coal extraction at the No. 8105 coalface, under the influence of coal extraction, there were some variations in stress distribution of the surrounding rock in the road- way. The stress concentration occurred in coal and rock body in front of the coalface and in the surrounding rock of the roadway, resulting in the deformation of surrounding rocks in the roadway. The roadway surface displacement was monitored in the No. 5105 roadway within a 100 m range in front of the No. 8105 coalface, the monitoring results are shown in Fig.72b. As can be seen from the figure, there was no change for the roadway surface dis- placement exceeding 78 m from the coalface, the roadway deformation affected by the coalface abutment pressure appeared from a location 78 m in front of the coalface. Within 78–66 m range in front of the coalface, large deformations occurred at the same time in the roof and two
walls. Subsequently, the deformation increased as the coalface advanced slowly. Within 36 m range from coal- face, with the advance of the coalface, the deformation began to dramatically increase. In the location of the monitoring station 5 m away from the coalface, the two side walls of the roadway moved closer by 280 mm, the roof and floor closer by 210 mm. However, in general there was no obvious damage to the surrounding rock of the roadway during the coal extraction period, the amount of deformation was small, and the transportation and venti- lation requirements was fully able to be met.
7.4 Analysis on gas control
On the 9th of November 2010, the No. 8105 coalface and roof drainage roadway were connected. After the connec- tion, a 2BEC80 gas drainage pump was used to drain the gas, the gas concentrations were maintained between 1.5 %
Fig. 73 Roadway support at the No. 5105 roadway
Fig. 74 Monitoring curve of absolute gas emission volume
Fig. 75 Variation curves of gas concentration at the upper corner of
and 2.2 % in the high drainage roadway, between 0.2 % and 0.3 % at the upper corner of working face, between 0.2 % and 0.3 % at the tail of the rear scraper conveyor, and between 0.15 % and 0.25 % in the return air. Thus, the gas management at the coalface achieved very good results.
The gas emissions at the coalface were monitored before the coalface extraction, from coalface extraction to the initial caving of the upper roof, to first gas drainage at the upper corner of coalface, during the gas drainage at the upper corner of coalface, and from the gas drainage at the upper corner of the coalface to the gas drainage in return airway and high drainage roadway. The monitoring curve of absolute gas emission volumes are shown in Fig.74.
Figure75shows the gas concentration variations at the upper corner of the coalface, the coalface and in the return air since the extraction of the No. 8105 coalface. As can be seen from the diagram, coalface gas emission rates grad- ually increased, the volume of gas drainage increased, while the gas concentration at the upper corner of the coalface, in the coalface and the return air gradually leveled off, and achieved a zero gas overlimit.
Site monitoring results showed that: the technology of roof high roadway drainage combined with other drainage can significantly reduce the gas concentration at the return air corner of the goaf, control the gas emission from the goaf to the coalface, play a role of ‘‘split-flow’’ of gas flow, and reduce the gas concentration in the goaf. The effect for preventing the gas accumulation at the upper corner of the coalface was significant. The drainage rate in the goaf reached 40 % of total gas emission in the goaf, achieving a zero overlimit, and ensuring a safe and smooth face advance. The gas control achieved a satisfactory result.
7.5 Analysis on fire prevention using nitrogen injection
Based on air quantity tests at different measuring points in the coalface, the amount of air leakage in the goaf was determined. With the installation of a bundle tube in the goaf and the connection of the bundle tube in the goaf with the main bundle tube monitoring system in the mine, the gas in the goaf was analyzed automatically. In addition, the airflow in the goaf was simulated in 3D to analyses the oxygen concentration in the goaf. Thus, a ‘‘three-zone’’ distribution was found in the goaf. Based on the finding, the nitrogen injection method and the amount of nitrogen injection were given at different face advance speeds.
Based on the actual mining conditions at the No. 8105 LLTCC coalface in Tashan mine, the nitrogen injection parameters were optimized, and the relevant nitrogen injection process and methods were developed. After cal- culation, an additional 3890 m3/h nitrogen injection was
required in addition to original 2500 m3/h, to enable oxy- gen concentration at the spontaneous combustion zone to drop to an average of 7 %. Pipe nitrogen injection tech- nology was used for fire prevention. The first nitrogen injection pipe along the goaf at the intake side of the coalface was buried. When the pipe was buried at a certain depth into the goaf, nitrogen injection commenced. Sub- sequently, a second nitrogen pipe was buried into the goaf (the shift interval of the nitrogen injection nozzle was 50 m). When the second nitrogen injection pipe was buried and started nitrogen injection into the spontaneous com- bustion zone in the goaf, the first pipeline of nitrogen injection stopped, and a new pipe was buried again for nitrogen injection. The process was repeated until the coalface extraction was completed.
The nitrogen injection method was determined based on fire prediction, in the case of nitrogen injection quantity of 2500 m3/h, and coalface advance speed of 1.4 m/d, a continuous nitrogen injection method must be used. When the advance speed is less than 1.4 m/d or if the coalface has stopped, the amount of nitrogen injection must be increased. When the coalface has stopped working for more than 68 days, the nitrogen injection quantity should be not less than 6390 m3/h; when the face advance speed is greater than 1.4 m/d, it would be appropriate to reduce the amount of nitrogen injection. Using the methods presented above, the occurrence of fire in LLTCC coalface can be prevented to achieve safe mining operation.